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By R J. Hardy

The aim of this study is to examine the volumetric and mass movement characteristics of load and haul operations using statistical modelling of relatively large samples of bucket loads that are transported as small samples in the form of truck payloads. Generally, the factors influencing bucket load variability are random, difficult to predict and to control. In the process of loading trucks the underlying variability of individual bucket loads will determine the variability of truck payloads. Bucket loads of material prepared by drilling and blasting for loading into trucks are fundamental quantities in open pit mine production. A study of payload data from a large surface mining operation indicates that bucket loads and truck payloads that are a small sample of bucket loads are normally distributed. Truck payloads have a reduced dispersion compared with bucket loads that is an inverse 0.5 power function of the number of passes per truck payload. Comparison of statistical modelling results with the æ10/10/20Æ policy promoted by Caterpillar Global Mining Inc, indicate that compliance with the æ10/10Æ part of the policy, applying to truck overloading is facilitated, as the coefficient of variation of truck payloads decreases, as a result of the increasing number of loading passes. Compliance with the more stringent æ20Æ part of the policy, relating to safe steering and braking is also facilitated, albeit at lower coefficient of variation for similar reasons.

Sep 26, 2011

By Vishal Shukla, Ashok Kumar, C. Raghu Kumar, D. P. Chakraborty

"Traditional iron ore beneficiation plants using washing, screening, and jigging processes require large volumes of process water. Generated slime during processing, its disposal and impact on water usage, is an important concern for any mining company. The challenge today, during extracting iron ore, is how to dispose of slime such that it is contained and stable, while maximizing water reuse and minimizing surface footprint. The slime disposal methods are sensitive to the environment. Two major technologies presently being practiced for disposing the slime across the world are pressure filtration and paste thickening. In this perspective an attempt has been made to do comparative study between paste thickening and pressure filtration for selection of suitable slime disposable technology for Noamundi processing plant, Tata Steel. From the studies, it is evident that by adopting pressure filtration technique it was possible to produce filtered slime, whereas filtration, rate reduces significantly with the presence of finer fraction. Further , the greatest advantage of paste thickening is volume reduction of slime. By considering the technology limitation and comparative lab test work result, it is opined that combination of filtration and paste thickening technology is favorable in terms of capital and operating cost and slime >45 µ should be filtered and < 45 µ slime should go for paste thickening. The study and the results are presented in details.INTRODUCTION The beneficiation processes always end up with an enriched portion of desired mineral - “Concentrate” and an enriched portion of undesired minerals - “Tailing”. The proportion and nature of tailings generated in every beneficiation process differ from ore to ore and reflects the characteristics of the feed processed. The traditional method of tailing disposal in many mineral processing industries has been to dispose them as thickener underflow at around 35% solids. Though, it had certain advantages of transportation with ease but contribute to a greater volume occupied in the tailing dam. This sustained practice and failure to utilize the space available in the tailing dam judiciously endangered the mining industry. The tailing management has become the toughest task to the mining industry and most of the mines are running out of space for disposing the fresh tailings as many of these are at the verge of getting filled."

Jan 1, 2018

By X. M. Chen

The selection of leach reagent mainly according to the ratio of useful mineral to gangue in mineral raw material and the structure of mineral raw material. Besides that the price of reagent, the capability of reagent decomposing mineral and its causticity on equipment is also supposed to be in regard. On common conditions, the scandium content in mineral raw material is low, even Sc concentrate of specimen-Yunnan rare earth ore which contain Sc used in this experiment only is 148 g/t. Moreover, in enlarging scale experiments the content of scandium in leach liquid is only 3.6g/t, meanwhile various and complicated impurity existing in it. At present, both domestic and abroad has succeed in enriching scandium from vitriol solution, while enriching scandium from hydrochloric acid only has been studied in some minority units, which is not reported to be practical in industry. Therefore, this item offer a new research issue on Sc?s extract, which is as well as the task of this research. Spectrum analysis, chemic component analysis and scandium content have been made in main mineral measurement on rare earth with Sc in the paper. Based on characters of sample, select hydrochloric acid, vitriol and nitric acid as leach reagent and make different blueprints to accomplish the experimental research about selecting leach reagent. In the aspect of rare earth leach scandium leach density is researched, leach liquid to solid rate, leach temperature, leach time, leach granule rate and leach solution. This experiment claims that in the optimal experimental ambience, leach rate of vitriol is 58.18%, the leach rate of nitric acid is 50.73% and the leach rate of hydrochloric acid is 59.34%. Among them, hydrochloric acid has the best effect, therefore this experiment choose it as the leach reagent.

Jan 1, 2014

By G. R. Sawyer

Although backfilling has been in operation for many years, the methods for selecting appropriate slurry dewatering aids have, in some cases, been rather unsophisticated. This has led to many misconceptions regarding the role of the dewatering aid and its various means of application. Testing procedures employed in the correct selection of any practically applied dewatering aid must mirror the actual working processes in which they are to be used. This constraint necessitates the development of complex testing methods when applied to backfilling. The current testing procedures were scrutinized and their various merits noted. From this study a series of testing methods was evolved which incorporate the best aspects of these testing procedures with novel ideas which would eventually result in the correct selection of the dewatering aid. A major consideration in the application of any of these methods is the modus operandi of the backfilling process, and the relevance of the tests to the process. This work attempts to consolidate the testwork performed in the past and suggest new approaches in the correct selection of acceptable dewatering aids for backfilling.

Jan 1, 1988

By G. A. Hufford

The cutoff grade applied to a disseminated orebody mined by open pit affects both the waste ratio, hence the cost of producing ore, and the ore grade, hence the quantity of metal recovered. As such it has a significant impact not only on the annual profit or loss, but also on the life of a mining property. Cutoff Grade Strategies Some commonly used objectives in calculating cutoff grades are as follows: Minimize Unit Cost of Product (1) This is found by adding all costs associated with each possible cutoff such as mining, processing, and marketing. This is divided by the quantity of marketable product obtained at each cutoff grade, then finding the minimum. Because the market value of the product does not enter the calculation it has no effect on this strategy: the same cutoff is selected in good times and bad.

Jan 1, 1986

By D. W. Cooper

The AMAX of Canada Limited Kitsault mine is located approximately 150 kilometers, by air, north of Prince Rupert, British Columbia at the head end of A1 ice Arm, just east of the southern tip of the Alaskan Panhandle. It is presently only accessible by sea or air, although a road is being constructed by the company which will provide access to Terrace, a commercial centre with a population of about 15,000. The topography is typically coastal mountainous with peaks rising to 2,100 meters around the head end of the inlet. The concentrator is at an elevation of 515 meters with the mine at 630 meters. This open pit molybdenum mine was previously operated between 1968 and 1972 by British Columbia Molybdenum Limited, a subsidiary of Kennecott Copper Corporation. AMAX of Canada Limited has spent $194 million to refurbish and expand this operation.

Jan 1, 1982

By P. Rooney

In due course, new smelters will have to be installed in order to meet the increasing worldwide demand for copper. This paper presents the results of a study made for the purpose of identifying the most desirable countries in the world in which to locate a new smelter. Both sociopolitical and economic factors were taken into account. Criteria developed and methods used to assess each possible location are described. The study indicated that the most interesting potential locations are Canada, -Chile, Portugal and the USA for new copper smelting complexes.

Jan 1, 1983

By Robert Warner

WHEN a new mine is opened, and often when an operating mine must adapt itself to physical or economic changes, a mining system must be selected in complete detail. In the past the plan chosen was usually the product of the local staff alone, however limited the training and experi-ence of that staff might be, which often resulted in the introduction of methods that were familiar to the staff, but not so well suited to the con-ditions as others unknown to the local men. A new management, noting the difficulties with the existing layout, might make modifications or complete changes, so that many mines in their history have seen a whole series of mining systems, each somewhat laboriously grafted on to its predecessor. This should not be interpreted as a condemnation of the older staffs for sticking to the tried systems, such as open square-sets in the early days at the Homestake, because they knew it would work after a fashion, nor does the statement apply to the technically trained, and usually broadly experienced staffs of most present-day properties. The older miners did not have the variety of tools and methods, nor the present freedom of transfer of information as to what the other man is doing, and so did the best they could under the circumstances. Finally, the higher grade of the ores made the choice less vital. The ever-increasing need to mine low-grade orebodies, with a fixed or even lower price for the metal produced, has caused an increasing need for low-cost mining. In most new mines this has simplified the problem of selection by eliminating from consideration all of the high-cost systems, but it has also complicated the choice by the necessity for refinement in detail and a close adjustment of the plan selected to all the conditions at the property. In these low-cost problems any textbook division of mining systems as applied to "beds, veins or masses" avails little as a guide. It should be realized at the outset that to&apos; be successful under modern competitive conditions a system of mining must be worked out with elaborate detail, with careful time scheduling and control of the various elemental operations, and that after it is introduced it must be run with a large amount of technical direction.

Jan 1, 1933

By George J. Simchuk

The selection of a stripping method and subsequently, the stripping equipment, are two of the most important decisions confronting a project engineering staff during the initial feasibility study of a new coal stripping operation. The selected stripping units average from 60 to 70 percent of the total capital outlay for mine equipment and their selection will affect the type, size, and quantity of the attendant equipment, such as loading shovels, drills, haulage trucks, and bulldozers. This paper is intended to report on an in depth study by Utah&apos;s Technical Services Staff regarding the selection of stripping equipment for the Goonyella Coal Mine in Queensland, Australia. We will touch briefly, on the objectives considered in this selection; the various alternatives which were explored; the general assumptions that were followed throughout the engineering and design of an appropriate stripping system; and computations necessary to determine the amount and size of stripping equipment. INITIAL EXPLORATION Utah Development Company (presently 90% owned by Utah Construction & Mining Co. and 10% by Australian investors) commenced systematic geological studies of the Bowen Basin coal fields of Central Queensland in 1961. The Blackwater Mine as shown in Figure 1 has since been established in this area and is now producing 3.0 million tons per annum of washed coking coal. Geological reconnaissance was then extended to the northern part of the Bowen Basin and in 1963 Utah&apos;s coal prospecting area covered 2444 square miles of Central Queensland. The authority to prospect north of Blackwater has since been trimmed to 810 square miles by relinquishments to the State.

Jan 1, 1971

By P Hagan, S Campbell

A study was undertaken using various methodologies to determine the optimum loading and hauling configuration for different haul routes at Callide Mine, a coal mine located in central Queensland. The study involved development of an effective flat haul model that was used to estimate production rates in truck and shovel circuits that employ two excavators of differing capacities then Queuing Theory to construct a capacity constraint model for scheduling purposes to assist in truck allocation and finally a cost model to determine the most cost effective fleet configuration.The model was calibrated using haul road design and performance data of trucks and excavators currently in use at the mine. The study examined the effects of cycle time for different truck and excavator fleet configurations on production rate and costs.The study found haulage cost was directly related to cycle time while production rate reduced with cycle time. The study indicates the approach taken in combining a cost model with a capacity constraint model offers a simple and effective method of scheduling an optimum number of trucks in a haulage circuit.CITATION:Campbell, S and Hagan, P, 2012. Selection of an optimum truck and shovel fleet based on effective flat haul and capacity constraint modelling, in Proceedings Eighth Open Pit OperatorsÆ Conference , pp 19-26 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 18, 2012

By P J. Ellen

The selection of the ore haulage system is a key decision made in the development of an underground mine. The haulage system is one of the components that define a mineÆs ability to respond to changes in mining inventory and market conditions. The complexity of this decision for Telfer Deeps was affected by the realisation during the feasibility process that the true scale of the potential underground mineralisation at Telfer was still to be quantified. This paper provides an overview of the process used in the Telfer Project Feasibility Study to select an ore haulage system for Telfer Deeps. The purpose is to highlight the value of impartially approaching tasks of this type with a robust process. The process should be able to respond to issues identified as the task progresses towards the objective of finding the best solution for an ore haulage system for the mine. For Telfer Deeps, this process resulted in the selection of a haulage system built around a shaft. The systems considered during the feasibility are discussed and the components of the haulage system constructed for Telfer Deeps are described.

Jan 1, 2005

By Piyush Rai, Suresh Kumar Sharma, Punit Paurush

The present study has been conducted in surface limestone mine to select the controllable blasting design parameters affecting the peak particle velocity (PPV) for assessment of ground vibration. The study aims at using two prominent statistical tools, namely, principal component analysis (PCA) and stepwise selection and elimination (SSE) techniques for identifying the key controllable parameters affecting the PPV. For determining the correlation coefficient between blasting design parameters and PPV from 44 field scale trial blasts, multi-variate linear regression (MLR) technique was done. It was found that the PCA and SSE eliminated a huge number of controllable parameters to identify the most important parameters, affecting the PPV. The PCA eliminated 6 number of parameters while SSE eliminated 5 number of parameters, and the coefficients of determination (R2) obtained were 0.616 and 0.584 for PCA and SSE respectively. The predictor equations were evolved, and these equations were used to validate the PPV results for another set of 21 field scale trial blasts. The predictor equations have been found to be fairly accurate in predicting the PPV values. Further, the PCA technique provides very near prediction of PPV with high degree of correlation in comparison to SSE technique. The paper highlights the role of state-of-art statistical tools in selecting the blasting design parameters affecting the PPV in field-scale blasting.

Mar 1, 2021

An extensive study was carried out to evaluate the performance of Outokumpu OK50, OK38 and Maxwell MX14 cells at ComincoÆs Red Dog flotation concentrator in Alaska. Tests were conducted by varying operating conditions of the cells installed in different circuits in the plant. The hydrodynamic and gas dispersion properties were measured and the metallurgical performance of the cells was evaluated for each condition used in the tests. This paper highlights the importance of superficial gas velocity, bubble surface area flux, gas hold-up, power consumption, residence time distribution and froth flow behaviour in optimising cell performance. The findings of these tests provide valuable insights into selection of the best cell operating conditions to optimise circuit performance.

Jan 1, 2000

By K. H. March

In the present very highly competitive condition of the bituminous coal business, it is becoming increasingly apparent that coal mines which do not lend themselves to efficient mechanization cannot long survive. It would seem that those mining operations best fitted to survive the struggle for existence are burdened with the considerable capital and maintenance cast of carefully selected mechanical and electrical mining equipment. The ever changing conditions of coal mine operations require equipment which is adaptable to a variety of uses; and investigation of the mechanical and electrical equipment in most mines will show that the operating duty of much of this equipment is considerably different from that duty for which it was originally selected. And it would appear that one of the essential qualifications for a successful mine management is a marked ability to get satisfactory operating results from equipment which has been transferred from some other operation, rebuilt, rearranged, or in some way adapted to the job at hand. As a striking example of such mechanical ingenuity, many of us recall that the first longwall jigging conveyor used in Nova Scotia was built in Sydney Mines from odds and ends in the scrap pile at a total cost of less than $50.

Jan 1, 1930

By N. Aslin, R. Clark, G. Heferen

The methods of harvesting copper from stainless steel mother plates has been investigated from fundamental and operational experience perspectives. At least five different harvesting techniques have been identified that have different effects on product quality, current efficiency, current distribution, growth morphology and ease of stripping. Each method is examined from fundamental principles to understand what occurs during the harvesting process and why the particular effects are manifested. Experience from plants operating the harvesting method is provided to show that the fundamental understanding is relevant. Best practice harvesting method and harvesting management are described.

Jan 1, 2014

By G. J. Dickman, W. N. Thomas

The factors involved in selecting, a drive system for a grinding mill, including gearing arrangement, motor selection, and electric-supply system limitations are reviewed. Equipment costs are evaluated on a dollar per horsepower basis and starting performance is analyzed, the latter making use of a recently developed computer program. The present interest in large-diameter grinding mills and the rapidly rising cost of the drive in proportion to the cost of the mill establishes a need to explore the drive system in detail. The drive for a 7 to 8,000 hp mill will cost from $350 to $600,000 which represents nearly half the total cost of the mill package, and requires as much building space as the mill proper. Some of the drive components require a longer time to manufacture and the installation time and costs are equal to those for the mill. It is apparent that the drive has become a dominant factor in the overall mill design consideration. A new plant requires a tremendous investment in both time and money. With the high degree of competition in the mining industry, low capital cost is an exceedingly important factor in the development of a project. It is just as important, however, to insure that the equipment is going to perform satisfactorily for the anticipated life of the project without adding unplanned operating, maintenance, and replacement costs. This is the position in which the engineer or designer usually finds himself and is really one of his primary functionsestablishing a realistic compromise between cost and performance. Large primary grinding mills can represent a major portion of the cost of a plant and the drive system accounts for a substantial part of this cost. These facts justify a careful analysis of the drive requirements and all of the factors involved in ultimately arriving at the best system for the lowest cost. A number of excellent papers have been published on various aspects of mill drives. The authors have used information from these papers in making a study of a drive for a 7600-hp, 32 x 12-ft autogenous mill. This was a broad investigation, including mechanical and electrical design, layout, performance, and maintenance considerations, and overall economics leading to the selection of the optimum system for that mill. This paper was developed from the study and includes information applicable to large mills in general. For a successful and economical application, the drive must be considered as a system and all of the components analyzed to determine their effect on the system. This requires the participation of the mill designer, gear, and other mechanical component manufacturers, motor manufacturer, plant designer, and operator. A careful analysis of the mill torque requirements is first necessary to provide a basis for selection of the mechanical and electrical drive components. Because of manufacturing limitations on the size and capacity of gearing available for these low-speed applications, an evaluation of costs for various gear ratios and configurations in conjunction with motor speeds is required. Consultation with the power-supply company is necessary to determine limitations on motor-starting disturbances and to analyze operating costs on the basis of existing rate structures and low characteristics. Finally, consideration must be given to the drive arrangement and its affect on plant layout, space requirements, and location of auxiliary equipment for the selected process. There are presently three basic schemes for driving a mill. The first is a direct drive by a motor operating at mill speed with no gearing. The second is by a higher-speed motor through a trunnion-connected gear drive. The third is by a motor through the more conventional pinion and ring-gear with the option of an interposing gear box to permit greater motor speed than pinion shaft speed. [(Fig. 1.)] Several manufacturers have or are presently developing motors which drive the mill directly without the use of gears. One arrangement has the motor armature intimately surrounding the mill shell-discharge spout or extension of the trunnion journal. The motor stator is then supported independently and surrounds the shell or trunnion-mounted rotor. Another variation utilizes an independent mill-speed motor connected to the trunnion by a drive shaft and flexible couplings.

Jan 1, 1972

By K Dudley, D Readett

The design of entrainment control processes in copper SX plants has been studied by use of models based on achievable plant performance. A number of alternatives have been assessed for their suitability in minimising the transfer of deleterious species from PLS into the electrolyte. The position of the wash (scrub) stage in the circuit, either before or after the loaded organic tank, has been the subject of considerable debate. The modelling shows that at ænormalÆ entrainment levels, the performance of the circuits are very similar. However at high entrainment levels the selection of the wash after the loaded organic storage has significantly lower transfer of contaminants. When considering the inclusion of a loaded organic coalescer the results of the modelling again show that the best circuit has the loaded organic storage/coalescer before the wash stage. The overall control of EW contaminants includes both a wash stage and an EW bleed flow. The total water use for these two flows is shown to go through an absolute minimum. The minimum flow rate is not greatly dependant on the total entrainment from the SX settlers. This can be a useful operating strategy if the quantity of high quality water for make up is limited by the capacity of the water treatment plant (eg RO). However the minimum cost option is to maximise the flow of water to the wash stage and minimise the EW bleed stream. The main driver for this is the cost of the cobalt sulphate contained in the electrolyte. Some results from the latest circuit modifications at the Nifty and Girilambone Copper Projects are presented to confirm the benefits of using both coalescing systems and a wash stage. The EW bleed has been reduced to a level that is less than the loss internally in the SX stage to stage entrainment. There has been an occasional need to add chloride to the EW to maintain sufficient concentration to act as a growth modifier.

Jan 1, 2002

By Howard S. Avery

The wear resistance of an alloy in heavy mill liner sections may be quite different from that of the same alloy in small balls and other light sections. This fact was established by means of a well standardized laboratory test that was validated for ball mill service. The salient properties of mill liner alloys, including austenitic manganese steels, medium alloy hardenable steels, and abrasion resistance cast irons are described, critically compared, and quantitatively ranked in terms of mill liner service. Where they are tough enough the high-chromium and other martensitic cast irons (e.g. Ni-Hard) promise maximum economy, but their section sensitivity is a limitation. To overcome this an extensive research program studied over a hundred experimental alloys, ranking them for heavy section abrasion resistance by means of the laboratory test. Computer analysis of the results provided a parabolic formula in terms of composition, the apexes of the parabolas providing optimum wear resistance. From the extensive family of parabolas one apex was selected for intensive study in experimental liners, pilot plant production, development of quality controls, and finally full scale production. This is a 19% chromium iron critically alloyed with nickel and molybdenum and hardened by heat treatment. It has excellent resistance to various types of abrasion, but its chief advantage is in heavy mill liner sections, where it has about 25% better abrasion resistance than the best of the quality controlled chilled Ni-Hards. It is also tougher. After production problems were solved, a number of ball and rod mill tests were arranged on a guaranteed performance basis. Two of these, representing about 138,000 pounds of mill liner castings, have recently confirmed in service the superiority that was predicted by the laboratory tests.

Jan 1, 1969

By M K. Singh

Mining is nothing but an art - besides being a science - of extracting valuable minerals from the womb of the mother earth. In this gigantic task, a large number of machinery and equipment are employed judiciously. Wire ropes play a vital role in lifting or hoisting men and materials under hostile/adverse working conditions and constitute a major item of investment on machinery/equipment. The economy of using a wire rope is a concern for engineers, users and the planners. The economic use of wire ropes for mining purposes largely depends on the following factors;Mechanical design considerations - which include tensile strength of wires, elasticity, resistance to abrasion and crushing, fatigue values, static and dynamic loads on wire ropes, factor of safety, fleet angle, etc. Metallurgical consideration - Composition and heat treatment of wire materials, defects like inclusions and voids, surface quality factor, lubrications, micro structural parameters. Operational consideration - Seizing and cutting of rope, storage, care and maintenance, cutting back, etc. Environmental consideration - Temperature and humidity, pit condition, dust, corrosive atmosphere, etc. In the present paper, an attempt will be made to suggest steps for achieving optimum economy while using the wire ropes for extraction of precious minerals and coal keeping in view the points noted earlier. Once technical viability is established, economy is the ultimate consideration for an industry for safe hoisting of men and material.

Jan 1, 2000

By David J. Dunn

GENERAL NOTES Liner Design for Comminution Liners in crushers and mills are the mechanical link between machine and ore. The way liners transfer energy to media and ore determines to large degree the production rate, liner wear rate, maintenance cost, energy use, and availability of the comminution process. Plant capacity and mineral recovery economics, and return on investment, are thus strongly dependent upon these simple, low cost metal shapes of which so much is demanded. The complexity of dynamic relationships between liner configuration, material properties, ore characteristics, and operating parameters defy simplistic generalization. Rules-of-thumb and axioms of years past fall into disuse almost daily as new analytical methods, assisted by computers and sophisticated control systems, are applied to crushing and grinding processes. Recent literature is well supplied with references citing production benefits that improved liner designs produce. This chapter is intended to review especially promising liner designs and material developments and presents cameo case studies showing the benefits derived from their application to comminution circuits (Tables 1,2,4,5,7). Complete discussions of liner design and wear materials are beyond the scope of this chapter. Norman&apos;s comprehensive studies in references (1) and (2) and Mclvor&apos;s analyses of mill speed/configuration effects (3) are especially useful reading for designers and operators of crushing and grinding circuits. Changing trends in ore, product value, and costs of wear materials, interest, labor, and energy, make ongoing evolution of liner design equally important for existing mines and for projects still on the drafter&apos;s board. Original designs supplied as new equipment must, of necessity, be best estimates. These can usually be improved as operating experience replaces engineering forecasts. Advent of new materials or design concepts such as the Angular Spiral Liners (Korpi (4) ) or Norman&apos;s Bimetallic design (2) offer improvement potential that must be evaluated if a given circuit is to be kept competitive versus others making similar products.

Jan 1, 1982